Metallurgical process



June 19, 1951 E. R. GILLILAND METALLURGICAL PRocEss 5 Sheets-Sheet 1Filed July 50, 1948 .m n @Wm m ra .In a e dm es uc GH we sd d GM /2/(Reducing Gas hier June 19, 1951 E. R. GILLILAND 2,557,650

METALLURGICAL PROCESS I Filed July so, 1948 3 Sheets-Sheet 3 LQ///////// f//yy l AFiat-5 Gas from /ilezng Section 4Q f4 E?! Heamg Gas///f// /fw//Z/ 26 ///Gas Riss@gema-7xm Sezlmg .Section FIG.- 6

Patented June 19, 1951 UNITED STATES 2,557,650 METALLURGICAL PRocEssEdwin R. Gilliland, Arlington, Mass., assignor to Standard OilDevelopment Company, a corporation of Delaware Application July 30,1948, Serial No. 41,667

22 Claims.

The present invention relates to a process and apparatus for reducingmetal ores and purifying the reduced metal by sineltng. Morespecilically, the invention refers to improved means by which orereduction and metal srnelting may be carried out continuously in asingle treating stage using a single integrated metallurgical unit. Theinvention is particularly applicable to the oXidic ores of those metalswhich 'are readily oXidiZable at elevated temperatures, such as iron.

In the copending Gilliland application, Serial No. 751,760 filed June 2,1947, now Patent No. 2,526,472, an open hearth furnace type process andapparatus for the smelting of iron-type metals are disclosed, in whichpowdered impure metal is introduced into the molten slag forming the toplayer of the furnace charge, to be melted therein, and the heat requiredfor melting the metal is supplied. by continuously 'circulating the slagthrough a separate heating zone', sufficient time being allowed for themolten metal tc settle out of the slag into the lower molten metal layerbefore the slag enters the heating zone. As shown in said copendingapplication, this procedure makes possible the melting of powderedimpure iron-type metals while at the Isame time permitting maximumutilization of fuel and control of the carbon content of the metal,independently of the combustion required to supply the necessary heat. H

It has now been found that the same principle may be applied to thesubstantially simultaneous reduction and melting of oXidic iron-typemetal ores, in such a manner that purified metal Aof 'the pig iron orsteel type may be obtained directly from the oxidic ore in a singleintegrated unit.

In accordance with the present invention, the oxidic metal ore and thereducing agent are introduced into the molten slag layer of an openhearth-type furnace in such a manner that the metal oxide dissolves inthe slag and reduction to molten metal takes place in the slag layer.The molten metal is permitted to settle cut of the upper slag layer intothe lower molten metal layer. Thereafter the slag is circulated througha heating zone wherein it picks up the heat re'- quired for reductionand melting, whereupon the reheated slag is again contacted with themetal ore and reducing agent.

In this manner no separate reduction step is required, the same fuel maybe used for heating and reducing the ore, low grade ores and reducingagents may be employed, and ore may be converted into finished steel ina single operation.

The ore may be supplied in its natural state o r in a partially reducedform. For example, FezOa, FesOr or EeO, or mixtures of these'oxid'es maybe used, preferably in a iin'ely divided form.

Suitable reducing agents include such conventional reducing gases ashydrogen, carbon monoxide, gaseous or vaporous hydrocarbons, etc., aswell as solid or semi-solid materials, such as oil residues, coal, peat,lignite, etc., or coke made from these materials.

When using a reducing gas it may be introduced into the slag eitherthrough the underlying layer of molten metal or directly withoutcontacting the molten metal layer. These two methods are substantiallyequivalent if hydrogen is the reducing gas. However, when a carbonaceousgas, such as CO or a hydrocarbon gas, is used the effects obtainedydiffer substantially.

Introduction of a carbonaceous reducing gas through the molten metallayer increases the carbon content of the metal by a reaction of themolten metal with the gas. As a result, the melting point of the metalis reduced, permitting lower melting temperatures and fuel savings, thereduction of the'ore in the slag layer is assisted by the contact of theslag layer with the carbon dissolved in the metal layer, and thechemical reactivity of the residual reducing gas remaining after contactwith the molten metal is increased. On the other hand, the metalrecovered from the molten metal layer is of the type of pig iron ratherthan of steel.

None of these eifects is obtained when the carbonaceous reducing gas iscontacted directly with the slag layer and this method will be moreuseful whenever any one or more of these effects appear undesirable. Inthis case, it is preferred to contact the reducing gas with the slagimmediately upon the latters return from the heating zone because theslag temperature is highest at this point and the higher the temperaturethe higher is the rate of reduction.

The gas remaining after contact with the slag layer and reduction of themetal ore normally still contains substantial proportions of combustibleconstituents at a high temperature. In

accordance with a preferred embodiment of the invention, this residualreducing gas may be used to transport the slag, after completion of orereduction, metal melting and molten metal settling, from the maintreating zone of the furnace toy the heating Zone to serve therein as afuel for heat generation, as will appear more clearly hereinafter. Otheruses of this residual gas include preheating and/or preliminary partialreduction of the incoming ore, ypreheating of other solid feedmaterials, etc.

Solid reducing agents may be introduced directly into the slag layer inthe form of powder or lumps. Since these solids are normally lighterthan the slag, special provisions should be made to force the solidsthrough the molten slag as will appear' hereinafter. The solid reducingagent need not have a high mechanical 3 strength such as that ofmetallurgical coke because it does not have to support a high burden.

The metal oxide may be added together with the solid reducing agent inthe form of suitable mixtures or briquettes, or the metal oxide may beadded separately from the solid reducing agent. The first method has theadvantage that some reduction may take place even before the ore reachesthe slag. However, a careful control of the ratio of oxide to reducingagent is required in this case. If the oxide is in excess, oxide will belost with the slag removed from the furnace. If reducing agent is inexcess it will build up and slow down the operation. It is preferred,therefore, to combine both methods in such a manner that a small excessof reducing agent is supplied together with the ore in the form of amixture or briquettes and additional ore is added separately to the slagat a suitable point so as to control the oxide concentration of the slagat a desired level. Additional oxide may also be added to the meltingand settling zone intermediate between reducing and heating zones inorder to adjust the carbon content of the molten metal.

The gases resulting from the reaction between ore and solid reducingagents may be utilized in substantially the same ways as outlined beforein connection with the reacted reducing gases.

Slagging ingredients such as iimestone or the like may be added atvarious points of the system. Addition to the heating zone is preferredbecause the necessary preheat and heat of decomposition may thus beapplied directly rather than by slag circulation and the CO2 evolvedfrom raw slagging materials may be removed with the flue gases from theheating zone and thus prevented from consuming valuable rcducing agentand heat in the main treating zone of the furnace. When added to themain treating zone, the evolution of CO2 will aid in the stirring andmixing of the materials to be treated and/ or in a reduction of thecarbon content of the molten metal, depending on the exact point ofaddition. In certain cases these desirable effects may compensatesomewhat for the loss of reducing agent and may maire this type ofaddition or a combination of this type with addition to the heatingsection appear attractive.

In accordance with a further embodiment of the invention, fresh ore maybe added to the heating zone. In this manner, sensible heat may besupplied directly to the ore by a combustion taking place in the heatingzone, whereby the amount of heat to be supplied by circulating slag maybe substantially reduced.

Further and more specific features and advantages of the invention willappear from the following detailed description of the accompanyingdrawing in which Figure 1 is a longitudinal horizontal section of afurnace according to the present invention, along the lines B-B ofFigure 2;

Figure 2 is a longitudinal vertical section of the same furnace alongthe lines A-A of Figure 1;

Figure 3 is a section similar to that of Figure 2, of a modied form offurnace according to the invention;

Figure 4 is a longitudinal vertical section of the furnace of Figure 1along lines C-C, illustrating a further modification of Figures 1-3;

Figure 5 is an enlarged longitudinal horizon- 4 tal section throughelement 30 of Figure 4; and

Figure 6 is a longitudinal vertical section of the type of Figures 2 and3, illustrating another modification of the invention.

Before describing the drawing in detail, it may be observed that noattempt is made in the drawing to supply details of constructionalfeatures which are common in open hearth furnaces. Wherever possible theconstruction of the furnace is represented schematically, the purposebeing merely to illustrate the features of the present invention. Asidefrom these features, the furnace may be assumed to be similar to theconventional open hearth furnace.

Referring to Figures 1 and 2, the numerals l, 2, 3, and 4 designate thefour walls of the furnace defining the bed 9. Arranged substantiallydown the middle of the furnace is an upright wall 6 which extends fromwall 4 more than half the length of the furnace to a point intermediatethe center of the furnace and wall 2, rising from the bed 9 of thefurnace preferably to its closed roof 2| and dividing the bed into twosections 1 and 8. In the bed S of the furnace at a location close towall 4, a row of reducing gas inlets I2 is provided which extends oversubstantially the entire width of section 1. A chamber '28 is providedbehind wall 4, this chamber being substantially coextensive with themain body of the furnace.

That portion of wall 4 which is in section 8 is provided with atransverse slot or opening 3B arranged so as to be in about the midportion of the slag layer whereby slag continuously runs from the slaglayer into the chamber 29. That portion of wall 3 which forms a wall ofchamber 29 may be provided with a horizontal row of ports 3l so locatedas to normally be below the liquid level in chamber 29. These nozzlesmay be on the same level as, or above or below, the slot 30. Thatportion of wall 4 which is in section 1 has a port 32 slightly below thenormal liquid level in section 1 and chamber 29 and the liquid level inchamber 29 will normally tend to be substantially the same as orsubstantially higher than that in section 1 of the furnace proper. Theend wall Il of chamber 2'9 is provided with an exhaust flue 20 whichleads gases away from chamber 29.

Closely adjacent wall 4, the roof 2 l of the furnace is provided with aninlet pipe 22 for powdered ore. Through this pipe additional quantitiesof slagging constituents, when required, may be introduced together withthe ore. Also alloying constituents such as carbon and other metals maybe introduced at this point. This pipe near its discharge end may beprovided with a cooling jacket (not shown) which may extend through thel'oof. At a point of roof 2l further removed from wall 4 is anadjustable flue 26 which may serve to draw off gases evolved during thereducing and smelting operation or which may be utilized to feed ininert or reducing gases so as to provide a protective blanket over themolten mass.

It will be understood that the furnace is equipped with suitablyarranged heating means, which constitute no part of the presentinvention, for bringing the furnace to reaction temperature. The furnaceis charged in the usual manner so as to establish a lower layer 21 ofmolten metal and an upper layer 28 of slag.

Referring now to Figure 3, the section shown therein is similar to thatof Figure 2 with the exception of the location of the reducing gasfeeding means l2, which is now arranged above and in spaced relationshipto the bed 9 of the furnace at an .elevation determined by a weir 34extending across the width of zone 1 of the furnace. The top of Weir 39-approximately coincides with the level of the molten metal layer '21 sothat the reducing gas introduced through feeding means l2 passes onlythrough slag layer 28.

Figure 4 is a vertical longitudinal section through zone 8 of thefurnace illustrating an embodiment of the invention which may be used incombination with either one of the modifications of zone 1 illustratedin Figures 2 and 3. Rear wall li of the furnace bed is provided in Zone8 with an elongated horizontal slot 39 connecting Zone 8 with chamber 29at such an elevation that the bottom lof slot 39 will be within and itstop just above the slag layer 29 when the furnace is in operation. A gasfeed line 39 arranged in the furnace roof 2| connects slot 39 with ablower or pump (not shown) for the supply of an oxidizing gas such asair. Drawoif lines 38 and 99 serve the withdrawal of molten metal andexcess slag, respectively.

The system illustrated by Figures 1, 2, and 3 may be operated asfollows. To start up the furnace it may be charged with metal and slagand heated up to operating temperature so as to form molten metal andslag layers as indicated in the drawing. Line 29 may be closed and areducing gas lbe admitted through feeding means I2 in the general mannerknown in the art of starting up Bessemer converters. Simultaneous- 1y,finely divided ore and any slagging ingredients needed may be chargedthrough line 22 by any suitable feeding means, such as a standpipe,lockhopper, star feeder, etc. (not shown). The ore is dissolved in slaglayer 28 and reduced therein to metal by contact with the reducing gas.The molten reduced metal settles through slag layer 28 into metal layer21.

As a result of the continuous withdrawal of metal and slag throughdrawoffs 38 and 99 from the end of zone 8 both layers 21 and 28 movecontinuously in the direction of the arrow.

A portion of the slag is passed through port 39 into chamber 29 whereinit is heated and pumped as described in the copending applicationidentified above. The heated slag in 29 flows rthrough port 32 intosection 1 and supplies the necessary heat for the reactions taking placein sections 1 and .8 of the furnace. The spent reducing gases withdrawnthrough port '29 may be used as heating gases for section 29 and forthis purpose introduced through ports 3l.

Referring now to Figure 4, port 32 is covered by liquid and line 26 isclosed, the gas remaining after reduction and still containingsubstantial proportions of combustible constituents is likewise forcedin the direction of the arrow and through slot 39 into chamber 29. Thevelocity of this residual gas may be readily so controlled that arelatively strong flow of at least the upper strata of layer 2B takesplace beyond drawoff t9 through channel 39 into chamber 29, asillustrated in greater detail in Figure 5 of the drawing. The relativelynarrow cross-section of channel 39 aids in establishing a pumping actionof the gas resulting in a strong current of slag from zone a intochamber 29. Depending on the gas velocity through, and the design of,port 30, this effect may be due to gas-liquid friction, slugging action,or spraying action, or to a combination of these causes.

Preferably preheated air admitted through pipe 3 6 `is admixed with thegas in channel 3.0 to ,form a ,combustion mixture which burns in channel30 or chamber 29 and heats the slag 'therein to the desired temperature.Flue gases are withdrawn through exhaust 2U.

It is preferred to maintain in sections 1 and 8 a pressure substantiallyhigher than that in chamber 29. This may be accomplished either byfeeding the reducing gas at an increased pressure or by applying suctionto exhaust 20, for example by connecting exhaust 29 to a suitable stack(not shown) or by any combination of these means. As a result of thispressure difference, the slag level in chamber 29 is higher than inzones 1 and 8 and the gases flowing through channel 39 bubble through alayer of liquid slag into chamber 29. In order to insure propercirculation of slag from the low pressure chamber 29 into high pressurezone 1, port 32 is preferably arranged at a lower level than slot 39, sothat the liquid column above port 32 in chamber 29 is higher than thatabove slot 39. If desired, additional combustion gases, such as suitablemixtures of gaseous fuel with air, may be introduced through ports 3linto chamber 29.

For a better understanding of the embodiment of the inventionillustrated in Figures 1 5, reference is made to the following operatingdetails for a typical operation in a furnace of the type described withreference to Figure 4, in which iron ore is simultaneously reduced andmelted.

Example I I. Daily charge;

A. Iron ore (52% Fe 46 tons B. Natural gas-1,350,090 cubic (60 F.-latm.) CE1-98% N2-2% C. Limestone- 25.3 tons (added to section 29) II.Daily production:

Pig iron-22.3 tons Operating temperature--2600 F. Air preheat-l300 F.Operating pressure- Section 29, l atmosphere absolute Section 8, 16inches H2O gauge The embodiment shown in Figure 6 is adapted to the useof a solid reducingagent in place of the reducing gas used in thesystems of Figures 1 5. Figure 6, in which apparatus elements similar tothose appearing in the preceding figures are identified by the samereference numerals, is a longitudinal vertical section through zone 1,similar to Figures 2 and 5, but modied to suit the special purpose ofusing solid reducing agents.

The construction of the furnace section illustrated by Figure 6 issubstantially the same as that of Figures 2 and 3 with the followingexceptions. Gas feeding means i2 are omitted. A bridge wall 52 extendsfrom roof 2l downwardly to a point well within the intended slag layer28, but above the level of metal layer 21, across the entire width ofzone 1. Bridge wall 42 is arranged in close proximity to solids feedline 22. A gas passageway Il@ penetrates bridge wall d2 in the directionof the longitudinal furnace axis at an elevation closely above the levelof slag layer 28. Line 46 has the form of a solids feed line rather thanthat of a gas pipe.

In operation, the furnace may be started up substantially as outlinedbefore. VWhen the feet proper temperature is reached, ore and solidreducing agent, such as coal, coke, peat, lignite, etc., are fed throughline 22 by suitable feeding means as outlined before. rhe solids may befed either separately, or in the form of preformed mixtures orbriquettes. slagging ingredients may likewise be supplied through line22.

Since the solid reducing agent is normally lighter than the slag itwould swim on top of the slag layer and the contact with ore dissolvingin the slag would be insufficient for an eicient reduction to takeplace. This diiliculty is overcome by the arrangement of bridge wall 42.The slag and molten metal layers move continuously beneath bridge wall42 in the direction of the arrow, impelled by metal and excess slagwithdrawal through drawofis 38 and 40 and by the pumping action ofreaction gases passing through passageways 44 and 30 in the direction ofthe arrow. However, the solid reducing agent swimming on the slag isheld up by bridge wall 42 and allowed to build up to a substantialheight of, say, several feet above the slag level. As a result, aconsiderable portion of reducing agent is forced into the slag wherebyproper contact with the dissolved oxides is accomplished. rlhe solidreducing agent may even be forced all the way to the bottom of thefurnace so that the molten slag and metal oxide filter through a bed ofsolid reducing agent whereby excellent contact and reduction areaccomplished.

The gases evolved by the reduction reaction are forced throughpassageway 44 and pass through settling and reducing zones 'I and 8 andslot 30 into chamber 29, in the manner described in connection withFigures 1-5. The combustible constituents of the gas are burned inchamber 29 to supply at least a portion of the heat required by theprocess. Additional gas may be added to section 29 to supply additionalheat.

If desired, additional metal ore may be added to the main section ofzone 1 through line 46. In this manner, the oxide concentration of theslag and with it the carbon concentration of the metal, may becontrolled at desired levels.

The following operating details will serve to further illustrate theembodiment of the invention described with reference to Figure 6.

Eccample II I. Daily charge:

A. Iron ore (52% Fe)-51 tons B. Coal (non-coking sub-bituminous)- 75tons (1/2 added through 22, 1/2 used for producer gas employed insection 29) Total carbon, 55.5 wgt. per cent Hydrogen, 6.2 wgt. per centSulfur, 0.3 wgt. per cent Nitrogen, 0.8 wgt. per cent Oxygen, 33.9 wgt.per cent Ash, 3.3 wgt. per cent Fixed carbon, 43.1 wgt. per centVolatile matter, 29.3 wgt. per cent Moisture, 24.3 wgt. per cent Heatingvalue, 9,400 B. t. u./pound (as received basis) C. Limestone-26 tons(added to duct 22) II. Daily production:

Pig iron-25 tons Operating temperature-2700 to 2800o F. Operatingpressure- 1 atmosphere Aii` preheat-130W F.

It is noted that in all embodiments of the invention described above,the location of the reduction zone A should be in closest proximity tochamber 29 in order to insure most efficient heat utilization andhighest reduction rates. The gases leaving the reduction zone A may be,wholly or in part, passed up line 22 countercurrently to the solidscharged in order partially to reduce the ore and to preheat all solidscharged through line 22. Particularly in this case, it may be desirableto feed additional combustion gases through ports 3l as indicated inFigure 1, in order to supply the necessary amount of heat and to aid inthe proper circulation of the slag. Flue gases from line 20 may berecycled to the system for similar purposes.

Special provisions may be made for the addition of slag and/or metal oredirectly to chamber 29. In this manner, the heat required to preheat anddecompose the slag and to preheat the ore may be applied directly ratherthan by circulating slag, and the CO2 evolved by slag decomposition iskept away from the reducing zone. Substantial savings in reducing agentand fuel requirements may be secured by this modification. Freshslagging ingredients should bc added directly to zone l only if thisappears desirable for the control of the carbon content of the metaland/o1- because of the stirring action of the CO2 evolved.

It will be apparent that the method and apparatus illustrated areamenable to considerable change in detail without suffering any changein essential character. While the particular embodiments illustratedpossess many unique features of construction and arrangement of parts,it is possible to design a suitable apparatus entirely different inappearance and general organization from that illustrated while stillutilizing the principle of supplying heat to the reducing and smeltingoperation by circulating slag through a separate heating zone andreturning it to the furnace bed. Such changes in design and arrangementare contemplated within the scope of the present invention.

The foregoing description has been more specifically directed to thereduction and purification of iron ores. It will be appreciated that theprinciples underlying the method described are applicable to any casewhere an ore the metal of which is highly reactive with oxidizing gasesat elevated temperature is to be purified by smelting.

While the foregoing description and exemplary operations have served toillustrate specific applications and results, the invention is notlimited thereto. Only such limitations should be imposed on theinvention as are indicated in the appended claims.

What is claimed is:

l. A method for reducing oxidic ores of metals highly reactive withoxidizing gases at elevated temperatures, which comprises establishing alayer of said metal in molten form covered by a 'layer of molten slag,continuously withdrawing slag from said slag layer, heating saidwithdrawn slag by contact with combustion gas containing oxidizingconstituents in a separate zone, returning the heated slag to the moltenslag layer, so as to furnish the heat supply required for reducing andmelting said ores, continuously feeding oxidic metal ore to said slaglayer, continuously feeding a reducing agent to said molten slag layer,Whereby metal ore is reduced to molten metal in said molten slag layer,settling reduced molten metal from said slag layer into said metal layerprior to the passage of said withdrawn slag to said separate zone, andrecovering reduced molten metal from said metal layer.

2. The method of claim 1 in which gases remaining after the reactionbetween said ore and said reducing agent are burnt to supply heat tosaid separate zone.

3. The method of claim 1 in which said reducing agent is originally agas.

4. The method of claim 3 in which said reducing gas is supplied directlyto said slag layer.

5. The method of claim 1 in which said reducing agent is originally agas and said reducing gas is introduced into said metal layer and passedfrom said metal layer into said slag layer.

6. The method of claim 1 in which said reducing agent is originally acarbonaceous solid.

7 The method of claim 6 in which said solid is supplied in subdividedform to said slag layer.

8. The method of claim 7 in which said solid is supplied in such amountthat it builds up through a substantial portion of the depth of saidmolten slag layer and said molten slag layer is forced through saidbuilt up solid.

9. The method of claim 1 in which said reducing agent is originally ahydrocarbon gas and said hydrocarbon gas is introduced directly intosaid slag layer at a point in close proximity to the point at which saidheated slag is returned to said molten slag layer.

10. The method of claim 1 in which said reducing agent is originally acarbonaceous solid and said ore and solid are charged together in theform of preformed mixtures.

11. The method of claim 10 in which said mixtures contain excessreducing agent and additional ore is supplied to said molten slag layerafter reduction has taken place in said slag layer.

12. The method of claim 1 in which slagging ingredients are added tosaid separate zone.

13. The method of claim 1 in which fresh ore is added to said separatezone.

14. The method of claim 1 in which said ore is an oxidic iron ore.

15. A method for smelting oxidic iron ores, which comprises passing amolten slag layer through a circuitous path comprising a reducing zone,a settling zone and a heating zone, in the i order stated, supplyingsubdivided ore and a reducing agent to said slag layer in said reducingzone, maintaining in said reducing zone a temperature adapted to meltand promote reduction of said ore in said slag layer in said reducingzone, settling molten reduced iron from said slag layer on its paththrough said reducing and settling zones to form a lower molten ironlayer, recovering reduced molten iron from said settling zone,contacting said slag layer with hot combustion gases in said heatingzone to heat the slag so as to furnish the heat supply required forreducing and melting said ores, and returning slag so heated to saidreducing zone.

16. The method of claim 15 in which gases produced in said reducing zoneare mixed with oxidizing gas, passed together with said slag to saidheating zone, and burnt to generate heat required in said heating zone.

17. The method of claim 15 in which an extraneous fuel gas and anoxidizing gas are supplied to said heating zone to generate bycombustion heat required therein.

'18. The method of claim 15 in which the liquid level in said heatingzone is substantially higher than that of said settling zone.

19. The method of claim 15 in which the liquid level in said reducingzone is higher than that of at least a portion of said settling zone.

20. The method of claim 15 wherein a substantially higher pressure ismaintained in said reducing and settling zones than in said heatingzone.

21. The method of claim 15 in which gases produced in said reducing zoneare mixed with an oxidizing gas, said slag is pumped from said settlingzone to said heating zone through a coniined, extended narrow path witha hot mixture of produced and oxidizing gas and said hot mixture ofproduced and oxidizing gas is burned to generate heat required in saidheating zone.

22. In the method of melting metalliferous solids by passing liquid slagthrough a substantially horizontal circuitous path comprising a meltingzone, a settling zone and a heating zone and supplying said solids tosaid melting zone at a point adjacent to said heating zone, theimprovement which comprises pumping said slag from said settling zone tosaid heating zone through a confined extended narrow path with a streamci" hot gases.

EDWIN R. GILLILAND.

REFERENCES CITED The following references are of record in the le ofthis patent:

UNITED STATES PATENTS Number Name Date 453,227 Wilson June 2, 1891828,583 Thiel Aug. 14, 1906 894,383 Imbert July 27, 1908 914,622Wikstrom Mar. 9, 1909 1,031,490 Thomson July 2, 1912 1,160,822 BeckmanNov. 16, 1915 1,313,309 Mambourg Aug. 19, 1919 1,328,636 Loke et al Jan.20, 1920 1,500,651 Smith July 8, 1924 1,535,109 Davies Apr. 28, 19251,647,608 Corsalli Nov. 1, 1927 1,868,666 Langer July 26, 1932 2,051,463Barker Aug. 18, 1936

1. A METHOD FOR REDUCING OXIDIC ORES OF METALS HIGHLY REACTIVE WITHOXIDIZING GASES AT ELEVATED TEMPERATURE, WHICH COMPRISES ESTABLISHING ALAYER OF SAID METAL IN MOLTEN FORM COVERED BY A LAYER OF MOLTEN SLAG,CONTINUOUSLY WITHDRAWING SLAG FROM SAID SLAG LAYER, HEATING SAIDWITHDRAWN SLAG BY CONTACT WITH COMBUSTION GAS CONTAINING OXIDIZINGCONSTITUENTS IN A SEPARATE ZONE, RETURNING THE HEATED SLAG TO THE MOLTENSLAG LAYER, SO AS TO FURNISH THE HEAT SUPPLY REQUIRED FOR REDUCING